Zinc is mainly produced by a roast-leach-electrowinning process route, which accounts for about 80% of the world primary zinc production. This process produces high quality zinc with very high zinc recovery (&gt;95%) but suffers from the several major drawbacks: (1) it uses electricity as reductant, hence its operating cost is high; (2) it produces iron containing leach residue either as jarosite or goethite, whose disposal poses an increasing environmental problem; (3) the elimination of sulphur from concentrate takes the form of sulphur dioxide necessitating the production of sulphuric acid which finds an increasingly limited market; (4) the process is not amenable to the treatment of secondary zinc feed materials containing high impurity levels, such as electric arc furnace dusts from the steel industry and other complex sources.
The so-called Imperial Smelting Furnace (ISF) process is the major pyrometallurgical process to treat zinc-lead concentrate. The advantage of this process over the electrolytic zinc process is its flexibility in treating high-lead, bulk concentrates with high recoveries of lead, copper and precious metals. However, this is a two-stage process involving sintering of Zn-Pb concentrates followed by blast furnace reduction. The process suffers from the need for large amounts of high-quality coke and sinter, which contribute to a large portion of the capital cost and results in serious environmental problems.
Increasing efforts have been made in developing new pyrometallurgical processes that can solve the above mentioned problems associated with the existing processes, especially in Japan, Australia and Europe, where the cost of electric power is higher than in North America. Several options are available as processes for front-end smelting:
conventional roasting followed by melting/slag fuming; PA1 two-stage bath smelting: oxidation followed by reduction; PA1 Outokumpu--or Kivcet-type flash oxidation smelting, with fuming of the high zinc slag; PA1 direct fuming of zinc from sulphide ores and concentrate; PA1 direct smelting in a molten copper bath (another type of two stage process); PA1 difficulty of condensing zinc vapour in the presence of high concentrations of SO.sub.2 ; PA1 undesirable formation of sulphur dioxide and hence the necessity to capture this gas, usually as sulphuric acid; PA1 the inability to treat significant quantities of low grade zinc-iron oxide residues; and PA1 the requirement of multiple stages in the processes.
Except for the direct fuming of zinc from sulphide concentrates, most of the above processes comprise two stages. The two-stage oxidation-reduction processes presume ability to handle a very high zinc slag from the first stage, and this could be a major impediment.
U.S. Pat. No. 5,178,667 discloses a two-stage bath smelting process for making zinc and lead from a sulphide concentrate. In this process, an iron-silicate slag or iron-silicate slag containing lime is formed and the incombustible sulphide concentrate flux and oxygen enriched air are blown into the slag to cause a reaction. As a result of this reaction, the major part of the zinc and part of the lead in the sulphide concentrate are dissolved in the slag. A reducing agent such as heavy oil, pulverized coal, powdered coke, or the like is blown through the resulting slag, and the zinc and the lead in the slag are volatilized, and subsequently condensed to molten zinc and molten lead.
U.S. Pat. No. 5,131,944 discloses another two-stage bath smelting process for making zinc, in which zinc sulphide concentrate is smelted with oxygen to form high zinc slag which flows to the second part of the vessel where it is reduced and volatilized to the gas phase by injection of coal and oxygen; the zinc laden gas is then condensed in a lead splash condenser.
A similar bath smelting process was proposed in U.S. Pat. No. 4,514,221 in which zinc calcine produced from the fluid bed roasting of zinc sulphide concentrate, is injected with coke fines and oxygen into a slag bath where zinc is volatilized and subsequently condensed in a lead splash condenser. The results of a pilot test of 15 tonnes per day showed that the overall zinc recovery is 70% to 74% with a coke consumption of 1.4 to 1.5 tons per ton of metallic zinc produced.
Flash smelting of zinc calcine using coal and oxygen was originally proposed by Kellogg in Lead-Zinc-Tin '80, ed Cigan, Mackay and O'Keefe, TMS-AIME, 1979, 28-47, and pilot tested at the Outokumpu Research Centre, Pori, Finland as reported by Asteljoki et al. in Extractive Metallurgy, 1989, 3-27. With a coal to calcine ratio of 1.1 to 1.3, they have achieved 89% to 93% zinc volatilization to the gas phase (zinc condensation was not tried during the tests).
A one-step smelting method for zinc sulphide concentrates using air and fossil fuel to produce zinc vapour containing gas which is then condensed in a splash condenser has been proposed by Yazawa in Metall. Trans. B., 1979, 10, 307-321, Davey in Australian Patent No. 59505/90, and Davey and Turnbull in Proc. of Australia/Japan Extractive Metall. Symp., Sydney, Australia, AIMM, July 1990, 23-29. According to these authors, a gas containing about 6% zinc could be produced by flash smelting zinc concentrates at about 1300.degree. C. with pulverized coal or other fossil fuel and air; this zinc vapour, in an SO.sub.2 containing gas, would then be condensed in an ISF-type lead splash condenser to produce liquid zinc. Iron, small amount of zinc and other gangue materials will form discard slag with flux. Thermodynamic calculations made by these authors have shown that, under a well controlled oxygen potential (through the control of CO to CO.sub.2 ratio), about 95% of zinc in the feed could be volatilized. Because of the low equilibrium partial pressure of zinc at the process temperature, a diluent gas, such as nitrogen, must be introduced to achieve high zinc recoveries to the gas stream. This increases the size of the gas stream to be handled, thereby increasing energy requirements, and capital costs.
Though thermodynamically feasible, there are doubts regarding the practicality of the above proposed process. According to the calculation of Davey and Turnbull identified above, the recovery of zinc is very sensitive to the ratio of oxygen to zinc: a 10% deviation from the optimal value would result in a 40% to 60% decrease in zinc recovery.
Another major difficulty would be the condensation of the zinc from SO2 containing gas due to the tendency for back-reaction of zinc vapour with SO.sub.2 gas at lower temperatures.
WO 91/0831 proposes another one-stage bath smelting process for producing metallic zinc. In this process, zinc sulphide concentrate is fed into a slag bath in which a controlled ferric to ferrous ratio is maintained and reacts with ferric iron to form zinc vapour and SO.sub.2. Air or oxygen enriched air is injected into the slag bath to oxidize part of ferrous iron to ferric and to burn coal to provide the heat required by the process. The zinc laden gas generated from the slag bath is then shock chilled by a fluid bed heat exchanger to produce metallic zinc. Since the slag temperature is relatively low (1150.degree. to 1300.degree. C.), it is very difficult to control an optimum ferric to ferrous ratio at which zinc vapour is the major reaction product. The difficulty of condensing zinc in the presence of SO.sub.2 may also be one of the major problems.
GB 2,048,309 describes a method for recovering non-ferrous metal from a sulphide ore thereof. In this method, the ore is dissolved or melted into a molten sulphide carrier composition, such as a copper matte, which circulates in a metal extraction circuit. Thereafter the composition is contacted with oxygen, for instance in a converter, so that at least part of the ore is oxidized. The carrier composition absorbs the heat produced and transmits it to endothermic sites in the circuit. The metal to be extracted can be zinc or a molten sulphidic copper matte composition. The oxidation step converts the copper sulphide of the matte to copper which then is able to reduce the zinc sulphide ore directly into zinc.
It is characteristic of the method described in the previous paragraph that the process employs a reduced pressure vessel for recovery of the volatile materials as a metal or a sulphide thereof, or impurities by means of suction. Because the process is conducted at a reduced pressure, the process temperature is in the range of 1150.degree.-1350.degree. C. The heat required by the endothermic reactions in the contactor and the reduced pressure vessel is obtained by circulating an excessive amount of sulphide matte in the converter. The sulphide matte is heated in the converter or can further be heated with burners. In view of the above description, this process appears rather complex to carry out.
CA 2,096,665 describes a two-stage bath smelting process whereby zinc concentrate is fed along with metallic copper to an electrically heated vessel where copper reacts with zinc sulphide to form zinc vapour and copper matte. The matte is tapped from the primary vessel to a second smelting vessel operating under oxidizing conditions. In this converting operation, metallic copper is regenerated according to well known process chemistry. Waste elements, such as iron, form a disposable slag phase. The zinc laden gas from the primary vessel is collected in a zinc column condenser which allows partial separation of zinc from deleterious elements such as lead and cadmium. The process suffers from several drawbacks, for example (1) high cost due to the necessity of copper matte converting; (2) the formation of sulphur dioxide gas in the converting operation; and (3) requirement of a converter slag cleaning operation to recover copper losses in the slag.
Accordingly, the main drawbacks of the above mentioned pyrometallurgical processes for the extraction of zinc can be summarized as:
There is therefore a great need to develop a method for smelting complex zinc concentrates and residues which would allow the recovery of zinc of good quality and purity, while simultaneously avoiding any of the above disadvantages. Further, it would be preferable that such method be simple, low cost, and environmentally friendly.